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By John R. Turney

The success of the Mclaughlin acid pressure oxidatiun process has proved that a complex hydrometallurgical process can be an economically viable means of treating refractory gold ores. Homestake's Mclaughin acid pressure oxidation process was commissioned in August 1985 and ts currently operating at design tonnage of 3,000 tons per day. Mineralogical and process studies on the operating plant are discussed to show the improved understanding: of the process and its future applications. These studies are also compared with the initial process development work, thus illustrating the success of the initial scale up considerations.

Jan 1, 1989

By Michael J. Hannan, Thomas F. Torries

The nature of industry structure is examined and the structure of the aluminum, copper, lead and zinc industries are described by the use of factor analysis. Capacity is defined and the relationship between industry structure and capacity utilization is examined. It is demonstrated that excess capacity is possible in industries with competitive or monopolistic characteristics, but the effects of the overcapacity are vastly different in the two cases. In a competitive industry, cycles of overcapacity with attendant fluctuating prices are expected. A change in industry structure from monopolistic to competitive characteristics over time causes profits and economic rents to be transferred from the producer to consumers of the mineral.

Jan 1, 1989

By Russell B. Dix

Laboratory column leach tests are used routinely to evaluate gold and silver ores for heap leaching. The tests are fairly straightforward. The ore to be tested is placed into an appropriately sized leach column. Then an alkaline cyanide solution is applied to the ore and the effluent solutions are collected. The effluent solutions are then put through a column of activated carbon to recover the gold and silver in solution. Zinc precipitation may also be used for silver ores or if the ore contains a significant amount of soluble silver in addition to gold. The length of the leach time in the tests is mainly a function of how quickly the precious metals are recovered from the ore. At the completion of the tests, the tailings are assayed and the recoveries determined from the calculated head grades. The amount of ore to be used for a test is partly a function of its crushed size. The coarser the size, the larger the sample size for the column. A typical small-scale laboratory column test on ore crushed to -12 mm (-0.5 in.) would usually be conducted in a 100- to 150-mm-diam (4- to 6-in.-diam) column 2 m (6.6 ft) in height, containing about 20 to 45 kg (44 to 99 lb) of ore. A comparable test on the same ore crushed to 50 mm (2 in.) would be conducted in a 200- to 300-mm-diam (8- to 12-in.-diam) col¬umn 2 m (6.6 ft) in height, containing about 80 to 200 kg (176 to 441 lb) of ore. Larger-scale tests can be run by increasing the column heights and diameters. Conducting larger-scale tests results in increased costs for sample procurement, preparation, and running of the tests. The additional data obtained from

Jan 1, 1989

By R. B. Coleman, T. P. McNulty

HISTORICAL PERSPECTIVE Environmental protection is not a newcomer to the minerals industry, there having been numerous examples dating from decades ago. Sometimes the awareness of a need for preservation or restoration derived from the ethics and principles of the engineers and management responsible for development and operation of a property. On the other hand, pressures to clean up industrial messes frequently were applied by special interests such as farming and ranching, recreational and conservational groups, and local governing bodies. Mineral industry environmental controls in this century have therefore been a patchwork of good intentions and bad, clean operations and dirty ones. The Hoboken siphon converter for converting copper matte while effectively capturing sulfur dioxide at a relatively high concentration was developed in Olen, Belgium in 1928, a full sixty years ago. The alternative was closure of the Hoboken smelter by Belgian government health officials who had become intolerant of sulfur dioxide emissions which were causing serious respiratory ailments in the river valley in which Olen is situated. As early as 1958, the Anaconda Company had a full-time botanist on its metallurgical research staff with the responsibility of learning how to vegetate the extensive tailings deposits at Anaconda, Montana. By that time, considerable progress had already been made in restoration of Silver Bow Creek, which at one time was “dead” for as far as 40 miles down- stream from the point of entry of acid mine drainage in Butte.

Jan 1, 1989

By H. G. Naidu, Raj K. Singhai, S. P. Singh

Introduction The economics of open-pit mining de¬pend heavily on the cost of overburden removal. It is necessary to maintain a pit slope as steep as possible. Steeper slopes lower the volume of overburden to be handled, thus lower the cost of material handling. There is, however, a limit to which the pit slopes in any given situation can be steepened. Steepening beyond this point could jeopardize the pitwall stability, ren¬dering it unsafe. It might even result in curtailment of pit production. Several factors affect the stability of pitwalls. Some are natural, such as topogra¬phy and regional and local geology. Others are man made, such as pit geometry, ground water management, equipment, drilling and blasting, and the type of excavation. Greater control can be exercised in the latter with proper planning and by following a set of procedures based on field experiences. This article deals with one such control¬lable factor - drilling and blasting. The eco¬nomics of steeper pit slopes is discussed, followed by a record of drilling and blasting practices in western Canadian surface mines. Microprocessor-based instrumentation and its application to blasthole drilling is also described. Economics of steeper pit slopes The economics of steeper pit slopes is not a new issue. It should be reinforced, though, that pit slopes must be as steep as possible to ensure maximum recovery of minerals at a minimum stripping ratio. For operating, economic, and safety reasons, different pit slopes are maintained during different phases of pit development. Figure 1 shows a hypothetical ore body, cylindrical shape, with a radius of 225 m (740 ft) and located 30 m (100 ft) beneath the surface. The mining operations are planned at two different slope angles of 25° and 40°. The ultimate pit depth was established on the basis of the criterion of "break even pit." The variation of the stripping ratio at differ¬ent mine levels is shown in Fig. 2. Table I summarizes the eco¬nomics involved for two pit slopes, 25° and 40°. Figure 3 presents the profits during the life of the mine for these two pit slopes. This study shows that steeper slopes for an open pit operation are far more advanta¬geous than flatter slopes. But, what is the optimum slope at any one point in the life of a moving pit? And how can the stable slopes be maintained over the short- and long-term? Canadian blasting practices A survey of blasting practices in western Canadian surface coal mines revealed that, at most mines, hard consolidated overburden is drilled and blasted. Blastholes are drilled with rotary crawler-¬mounted rigs using bits in the 220 to 310 mm-diam (8.5 to 12 in.-diam) range. Typi¬cal hole depths are 12 to 18 m (40 to 60 ft) to accommodate the pit bench heights of 12 to 15 m (40 to 50 ft). For a 12-m (40-ft) high bench, a typical hole will be 14.6 m (48 ft) deep including the subgrade drilling of 2.6 m (8.5 ft). Half of this hole will typically be a powder column and half will be stemmed. For a 15-m (50-ft) bench, a typical hole will be up to 18 m (66 ft) deep including the subgrade drilling. Drilling patterns are usually 9 m (29.5 ft) spacing and 9 m (29.5 ft) burden. De¬pending on local geology and hole diameter selected, drilling patterns of 8 x 8 m, 8.5 x 7.3 m, and staggered patterns of 11 x 5.5 m

Jan 1, 1989

This paper describes the effects of physical variables in dewatering of mineral slurries with air as the filtrate displacing medium. Using a mathematical model with material related constants measured in laboratory dewatering curves are simulated for vacuum and pressure filtration. Four different mineral slurries are studied. Basic dewatering curves with variation in air-pressure are shown with iso-energy lines. Variables such as air pressure, feed concentration, cake thickness and cake formation pressure are discussed.

Jan 1, 1989

By Ronald J. McGregor

An overview or precious metals project development and start-up, which emphasizes the process decision. Logic which determines process flowsheet, project economics, and training required to train new operators for start-up. The process decision required to successfully develop a precious metals project requires a blending of test data, environmental considerations and operations experience. Decisions to use Merrill-Crowe or carbon adsorption/ electrowinning/replating, fixed pad or multi-use pad {in heap leach) are dependent not only on technology but also on ability to train new operators. The market for experienced operators is acutely depleted and almost all new operators will have to be "home grown". Experience from the Brewer project in South Carolina as well as other projects put together in the West will be discussed.

Jan 1, 1989

By M. J. Beus

A shaft is truly the "lifelinen of an underground mine. Damage to the shaft lining and guides as a result of ground movement can result in serious loss of production and extensive and continual repair. In some mining districts, annual shaft repair and maintenance costs resulting from high rock stresses and excessive displacement are in the millions of dollars per year. For many years, the Bureau of Mines has conducted research to characterize the rock mass and to develop structural guidelines to improve the design of accessways in deep mines. Much of this work has centered on determining the in situ conditions that affect the structural stability of shafts in the Coeur d'Alene Mining District of northern Idaho. Recent work by Beus and Board (1984) has concentrated on measuring the rock and support behavior during sinking of the Silver Shaft at the Lucky Friday Mine. These results are currently being used as a comparative data base for further basic research as well as to establish structural design criteria. The factors affecting the deign and structural stability of deep mine shafts are numerous. A design approach that use field data as input and directly compares field measurements with numerical models can validate the procedure and provide realistic design criteria. Naturally occurring "fixed? conditions such as magnitude, direction, and ratio of in situ stresses, geologic environment, and physical properties and constitutive relationships of the rock mass ere obviously basic to numerical modeling.

Jan 1, 1989

By Pollard D. M

The BHAS copper-lead matte leach process, developed in Port Pirie, South Australia, has been in operation on a commercial scale since 1984. This process is one of many developed in the past decade to recover copper via the hydrometallurgical route. In essence, the process involves a single stage agitative leach of 100% -75um matte in an acid chloride-sulphate lixiviant with gaseous oxygen as oxidant. The leach temperature is controlled at 85¦C to assist with reaction kinetics. After leaching, divalent copper is recovered from solution via a conventional solvent extraction/electrowinning circuit to yield cathode copper for sale The BHAS matte leach process has several unique characteristics cf. other commercial oxidative acid leach processes. These include i) the use of gaseous 02 as oxidant ii) H2SO4 addition as make up acid iii) Relatively low chloride ion addition (30 g/1) For this reason, extensive work was carried out into the optimization of leaching by Meadows (1985). Variables investigated included ò Oxygen flow-rate ò Agitation intensity ò Leach temperature ò Free acidity ò Chloride ion concentration ò Matte particle size The current work was conducted to determine the effect of cooling rate, and hence mineralogy, on the leaching characteristics.

Jan 1, 1989

By L. A. Melis, J. W. Rowson

"Cluff Mining has successfully operated a gold recovery plant at its Cluff Lake operation in northern Saskatchewan to recover gold f rom Phase 1 uranium leach tailings. The operating circuit consists of tailings repulping, grinding, cyanidation, carbon-inpulp, gold recov ery, and cyanide destruction by acidification, volatilization and re-neutralization. The process is described and operating data and metallurgical results from the operation are provided.INTRODUCTIONCluff Mining, a Saskatchewan partnership, is owned 20 per cent by Cameco and 80 per cent by Amok Ltd. /Ltée. (Amok), manager and operator of the partnership.The current operation includes the Phase II uranium mining and milling facilities, located at Cluff Lake, 700 km north of Saskatoon, capable of processing 300 000 tonnes o f ore to produce about 4 000 000 lb UPs (1500 tonnes of uranium) per year.The Phase I operation was initially established to process the very high grade ""0"" uranium orebody from 1980 to 1983. The leach tailings from the Phase I mill operation were found to contain elevated level s of gold . These high gold values were partially due to the gravity pre-concentration step included in the uranium recovery circuit, and to the naturally high levels of gold associated with the uranium mineralization in the""0"" orebody .During the Phase I operation some 6475 tonnes of leach tails were produced and stored in a total o f 2916 large concrete containers (vaults). This method of storage was, at the time, deemed to be the preferred method of temporary storage for the tailings.This temporary storage of high-grade gold tailings facilitated the reclamation and reprocessing of the leach tail s for gold recovery. The primary reason for reprocessing the Phase 1 leach tails was to provide for long-term permanent disposal of the material as an admixture with Phase II tailings in the Cluff Lake tailings disposal area. Although there was limited tonnage, the high gold content in the material made the in stallation of a gold recovery circuit economically justifiable."

Jan 1, 1989

By G. G. Marino

Little comprehensive information has been reported on the behavior of room-and-pillar mines. The objective of this paper is to present case data on mine failures in the Illinois basin for use in practice. Presented are results of an ongoing study and details on the site characteristics of cases where sags have developed on the surface. Site data are reported to show the geologic, mining, and sag conditions that existed. Sags mainly develop from pillar, floor, or pillar-floor failure. The character of the sags depends upon the type of mine failure as well as the overburden response.

Jan 1, 1989

By Bernhard J. Ralph

In refractory gold ores much of the gold occurs in sub-microscopic form in sulfide minerals, most commonly pyrite and arsenopyrite Bacterial leaching is being developed as a pretreatment for these minerals to aid the liberation of gold and a variety of extracting reagents. This paper reviews the nature of the principal micro-organisms involved, the effects of various physico-chemical factors on the rate and extent of extraction, and likely mechanisms involved.

Jan 1, 1989

By M. Timucin

Phase relations within the condensed Cu-Cu2S-PbS-Pb region of the Cu-Pb-S ternary were investigated. The boundaries of the ternary liquid-miscibility gap were redrawn, and the directions of the tie-lines were established by the quenching of samples equilibrated in evacuated and sealed silica capsules at 1473 K. The activities of lead in sulphur-saturated Cu-Pb alloys along the boundary of the miscibility gap were measured by the dew-point technique at 1473 K. From the tie-line distributions, measured lead activities, known activities of the Cu2S-PbS join, and the ternary Gibbs-Duhem integration, consistent sets of activities for lead and copper and partial pressures of sulphur were obtained along the ternary miscibility gap at 1473 K.

Jan 1, 1989

By T. O. Kuhl

The Royal-Mountain King Project is 7.2 kilometers (4.5 miles) northwest of Copperopolis, California. Reserves of 8,250,000 t (9,100,000 st) at 2.32 g/t (0.068 oz/st) gold will be processed at 2500 t/d (2750 st/d) from three pits. Production is scheduled to begin the first quarter of 1989. District geology consists of northwest trending sequences of Jurassic meta-sedimentary and meta-volcanic rocks. Three ore bodies are located along the Hodson fault which trends northwest and dips 40 degrees northeast. The Gold Knoll and North Pit ore bodies are hosted in carbonate and silica altered meta-volcanic rocks whereas the Skyrocket ore body is hosted in quartz stockwork zones in meta-sedimentary rocks. Gold is present as inclusions and micro-veinlets in pyrite.

Jan 1, 1989

By H. R. G. Steeghs

Recently developed organic binder systems have the advantage of eliminating bentonite contamination. This results in better chemically reactive taconite pellets. Both acid and fluxed pellets show increased R40 reducibilities, whereas mechanical and physical properties of the taconite pellets are not adversely affected. Blast furnace performance of these pellets is excellent, and substantial economic benefits are achieved.

Jan 1, 1989

By C. T. Christensen

The paper gives an outline of the design, installation and performance of a series of mooring and berthing dolphins for a major ferry berth outside the port of Elsinore, Denmark. The Dolphins are large cylindrical bottle shaped and open ended steel piles vibrated into the sandy seabed. They are subsequently equipped with prefabricated fender units. Vehicle access to the ferry is via a fixed ramp and a semi-submerged linkspan. Supplementary passenger access is via a covered gangway steel bridge from an existing gangway facility to a midship door in the ferry«

Jan 1, 1989

By R. E. Brewer

Pilot scale flotation tests were conducted with a 1-ft-diam column at the Barrick Resources (USA) Mercur mine, Tooele County, UT, to scavenge unleached (0.01 to 0.018 oz/st) gold values from the tailings of a carbon-in-leach (CIL) circuit. The optimal recovery was 35 pct of the unamenable-to-leaching gold, at a grade of 0.353 ounces of gold per short ton, using a mixture of dithiophosphate, xanthate, and diesel fuel as a collector. This was achieved by floating approximately 1.6 wt pct of the CIL tailings. The gold values are believed to be contained in arsenic-bearing minerals, attrited carbon, and carbonaceous material in the ore. Redox potentials were measured with a gold electrode during the conditioning stage prior to flotation. The correlation between flotation recovery and redox potential values could be used to control reagent additions. Treatment of a previously leached and weathered tailings material, averaging about 0.055 ounces of gold per short ton, was also investigated using the column. The tailings were first water leached at approximately 600 C, and the water-leached tailings, which contained approximately 0.03 ounces of gold per short ton, were then fed to the column flotation unit. The effects of pH and sodium sulfide addition on gold flotation recovery and redox potentials were examined. maximum gold recovery, 36 pct, was achieved at pH 8.5 with a sodium sulfide addition of 0.18 lb/st.

Jan 1, 1989

By A. W. Khair

This paper presents an analysis of fracture depth and intensity due to subsidence over the longwall panel. Sonic reflection techniques were utilized to determine fracture depth, and a variation of p-wave velocity was used as an indication of the fracture intensity. Instrumentation has been tested for consistency and accuracy first in the laboratory using small scale models, then applied in the field in two mine sites with different mining geometry and geologic conditions. A new sonic viewer was utilized for sonic velocity measurements. Regular hammering method was on acoustic source. The results were concurrent with monitored horizontal strain profiles and measured open fractures over the longwall panels.

Jan 1, 1989

By Gary Kahn

In this document the VODR is discussed from a technical point of view and is intended to give the reader a feel for the design of the instrument, its technical specifications, and its limitations. It is not intended as a tutorial on the use of the instrument in the field.

Jan 1, 1989

By Harold R. Newman

The performance of Ecuador's mineral industry continued to be affected not only by depressed oil prices in recent years but also by natural disasters. In March 1987, Ecuador canceled foreign debt payments for the rest of the year and appealed for international aid because of a series of earth-quakes that destroyed villages and crippled the nation's oil industry. The earthquake destroyed a key section of the country's main oil pipeline and forced a halt to production and shipment of Ecuador's most important export product. Ecuador produces 240,000 barrels of oil per day and exports 144,000 barrels daily. Oil accounts for 60% of the nation's hard currency earnings. The Government of Ecuador asked the Organization of Petroleum Exporting Countries and other oil-producing countries to assume Ecuador's oil sales commitments. Ecuador agreed to pay the countries back in an equivalent amount of oil when it was able to resume exports.

Jan 1, 1989